BG62180B1 - Chloride assisted hydrometallurgical extraction of nickel and cobalt from sulphide ores - Google Patents

Abstract

A process for the extraction of Ni/Co values from an ore orconcentrate comprises the steps of subjecting the ore orconcentrate to pressure oxidation in the presence of oxygen and anacidic solution containing halide, copper and sulphate ions toobtain a liquor containing Ni/Co values from the resultantpressure oxidation slurry. The liquor is subjected to a selectiveprecipitation treatment to obtain a solid containing Ni/Cohydroxide. The solid is subjected to a Ni/Co leaching stage withan ammonium solution to produce a leach solution containing Ni/Covalues and a residue. The Ni/Co values are separated by solventextraction to produce solutions suitable for electrowinning of Niand Co. The process also provides for the recovery of preciousmetals and other metals such as copper.

Description

The invention relates to hydrometallurgical treatment of ores and concentrates. In particular, it relates to the extraction of metals from sulphide and laterite ores in the presence of halide ions, such as chlorine ions.

BACKGROUND OF THE INVENTION

According to current practice, sulphide nickel ores are currently processed in various ways, in which the first step is almost always physical concentration by flotation to increase the nickel content in the concentrate, typically in the order of 0.5% to 2% to 7 to 25% nickel. Subsequent treatment of this concentrate is usually a pyrometallurgical treatment (fusion) to produce nickel matte or artificial high percent sulfide with about 20-70% nickel.

Then the matte is usually refined to nickel products using hydrometallurgical operations.

This combination of pyrometallurgical / hydrometallurgical processing of nickel concentrates is currently widespread in practice with some variants, in particular in the hydrometallurgy sector. Most of the processes extract some of the associated metals such as copper and cobalt. Additionally, the leach residue containing precious metals such as gold and silver, as well as platinum group elements, is often processed to further extract the precious metals contained.

There are some disadvantages to this processing scheme. The disadvantages associated with pyrometallurgical operations include:

(I) the production of furnace gases involves sulfur dioxide, which must be treated in an installation to produce sulfuric acid by-products, which can often be difficult to trade at close range. (The capital investment and operating costs of such acid production installations affect of the general economy of the process) (ii) losses of nickel and especially of cobalt in the slag product upon melting, often more than 50% of the incoming cobalt.

(iii) generally high melting costs, especially of low quality concentrates (<10% Ni); (iv) difficulty in treating certain concentrates with harmful elements: magnesium (Mg) and arsenic (As).

Hydrometallurgical operations for the treatment of nickel matte are significantly altered, with all known processes used having one or more of the following disadvantages:

(i) high costs for reagents such as sodium hydroxide or ammonia required for neutralization.

(ii) producing a large amount of by-product, such as ammonium sulfate or sodium sulfate, which are difficult to sell.

(iii) high energy costs due to the large temperature interval during the process.

(iv) a complex and costly technological scheme that results in high investment and high operating costs.

As an alternative to the pyrometallurgical / hydrometallurgical processes described above, a process is known that uses only hydrometallurgical operations so that concentrates are processed without melting. It uses a pressure leach with ammonia solution. This prevents most of the disadvantages associated with melting operations, unfortunately not all of the above disadvantages of known hydrometallurgical processes remain, and in fact are not even as effective overall as the best pyrometallurgical / hydrometallurgical processes.

Copper and nickel sulphide ores often also contain precious metals such as cobalt as well as precious metals such as gold and silver and platinum group metals. As these ores are usually of lower quality ores as far as in. . _____ _ copper and nickel and having a high sulfur to copper / nickel ratio, it is problematic to economically extract Cu, Ni and Co.

Some sulfide ores contain so little copper and nickel that the extraction of precious metals must be increased so that the process becomes economically viable. Due to the high content of pyrite in some ores, the extraction of gold through the conventional cyanide method is often difficult, which also makes it economically unprofitable to process the ore. The present invention provides a method for hydrometallurgical extraction of copper, nickel and cobalt as well as other metals from sulfide ores; also creates a method for the hydrometallurgical extraction of nickel and cobalt from laterite ores.

SUMMARY OF THE INVENTION

According to the invention, there is provided a method for extracting a non-copper metal from an ore or concentrate, comprising a step of oxidizing the ore or concentrate under pressure in the presence of oxygen and an acidic solution containing halogen ions and a source of bisulfate or sulfate ions to form a solution said non-copper metal, wherein said source of bisulfate or sulfate ions is selected from the group consisting of sulfuric acid and copper sulfate, which is hydrolyzed in said acid solution.

In addition, according to the invention, a method for extracting valuable nickel / cobalt from ore or concentrate is created, which consists of the steps of subjecting the ore or concentrate to oxidation under pressure in the presence of oxygen and an acid solution containing halogen, copper and sulfate ions for the preparation of a liquid containing valuable Ni / Co as a result of pulp oxidation; the liquid is subjected to selective precipitation treatment to obtain a solid phase containing Ni / Co hydroxide; the solid is leached with ammonia solution to obtain a leach solution containing valuable Ni / Co and waste.

The method may further include acid washing steps of the waste to obtain a washing solution containing the valuable Ni / Co and the waste removed and recycling of the washing solution for selective precipitation or optionally the washing solution is treated to recover the valuable Ni / Co.

Selective precipitation may include stages of liquid precipitation at pH 5-6 to precipitate iron, copper and zinc present in the solution; the resulting solution was precipitated at a pH of about 7 to 8 to form a solid phase containing Ni / Co-hydroxide.

The liquid containing the valuable Ni / Co may be obtained by subjecting the pulp to oxidation under pressure to neutralize at a predetermined pH, whereby the copper, iron and cobalt present in the pulp are in a solid phase and the valuable Ni / Co in solution or at optionally, a liquid containing valuable Ni / Co may be obtained, and the slurry is subjected to pressure oxidation to neutralize to a predetermined pH, wherein the valuable Ni / Co is in a solid phase and then the solid valuable Ni / Co is leached to acid. obtaining Ni / Co in solution.

The method may include steps of controlling the concentration of nickel in the leach solution to a maximum value of about 3 - 25 g / l, preferably 8 -10 g / l, and especially preferably 10 g / l.

The Ni / Co leaching step can be carried out with ammonium sulfate solution. The concentration of ammonium sulphate may be from about 150-250 g / l, preferably about 200 g / l.

The invention also provides a method of extracting valuable Ni / Co from a concentrate containing Ni / Co hydroxide, which includes the steps of subjecting the concentrate to Ni / Co leaching with an ammonia solution to obtain a leaching solution containing the valuable Ni / Co and waste; and controlling the concentration of nickel in the leach solution to a maximum of about 3 to 25 g / l.

The term "concentrate" in this application refers to any material in which the content of precious metal is increased to a higher weight percentage than the ore found in nature and includes artificially obtained sulfide ore, such as matte and precious metal sludge in the form of a solid phase such as hydroxides and sulfides.

The essence and advantages of the invention are explained by preferred embodiments of the invention described below.

BRIEF DESCRIPTION OF THE DRAWINGS The invention is illustrated with examples and explanations of the drawings, where:

Fig. 1 is a flow chart of a method for hydrometallurgical metal recovery according to the invention.

FIG. 2 is a flow chart illustrating more details of the solvent extraction process steps of FIG. 1.

Figs. 3A and B are a flow chart of another embodiment of the method of the invention for the extraction of precious metals.

FIG. 4 is a flow chart of a method for hydrometallurgical metal recovery according to another embodiment of the invention.

DETAILED DESCRIPTION OF THE PREFERRED EMBODIMENTS

The method according to the invention is suitable for the processing of copper ores, in particular sulphide copper ores, which also contain nickel and / or cobalt or nickel / cobalt sulphide ores with negligible copper content, as well as nickel / cobalt oxide / laterite / ores. In addition, nickel / cobalt ores containing other elements often considered as harmful, such as magnesium, arsenic and zinc, or elements that are valuable and valuable for extraction, such as precious metals gold and silver and metals of the group of platinum.

The ore or concentrate submitted for treatment may contain one or more sulfide minerals based on Cu.Ni, Co and Zn metals, often combined with iron And sometimes with other elements such as As, Sb, Hell and others.

Typical sulfide minerals based on the above metals are:

med chalcocite chalcopyrite nickel millerite pentlandite cobalt linait cobaltite zinc sphalerite marmatite

In this context, the metal: sulfur ratio is the ratio of the total base metals / Cu, Ni, Co / to sulfur in the concentrate, and this is an estimate of the quality of the concentrate.

. Typically, the metal: sulfur ratio ranges from 1.5 for high quality concentrates to below 0.2 for low quality concentrates. For concentrates that are predominantly Ni / Co, the metal: sulfur ratio is more often of a lower order, 0.2 to 0.8 / iron is explicitly excluded from these calculations, although it is present in almost all sulfide concentrates /.

The importance of the metal to sulfur ratio for the process lies in the fact that it influences the metallurgical changes in the initial operations of oxidation under pressure.

Various embodiments of the process according to the invention can be used to process a series of Ni / Co concentrates in which the metal: sulfur ratio changes from low to high, as briefly described above. However, in addition to the mentioned ratio, it is necessary to take into account another characteristic - the degree of oxidation of sulfur / to sulfate / during oxidation under pressure. The sulfur contained in the concentrate during oxidation under pressure is converted to either elemental sulfur (S7 / no sulfur oxidation / or oxidized to sulfate / SO ₄ '). Usually 70-95% of sulfur is non-oxidized and is obtained as elemental sulfur. In other words, the oxidation of sulfur (to sulphate) usually changes from 5 to 30%. It is considered useful to reduce the oxidation of sulfur; an important task of the present method is to achieve this. This is facilitated by the insertion of a sulfate or bisulfate source, such as H ₂ SO ₄ in the pressure oxidation step.

The importance of sulfide oxidation is that it produces an acid that can eventually be neutralized and it affects the distribution of copper and iron and other elements in the pulp resulting from oxidation under pressure. The high acid (low pH) slurry contains Cu in solution, while the low acidity (high pH) slurries contain solid copper as basic copper sulfate.

For low metal concentrates: sulfur and / or high sulfur oxidation, the flow chart shown in Figure 1 is a general case. This is referred to as mode C. During the oxidation under pressure 12, a sufficient amount of acid is obtained, so that it is necessary to neutralize this acid with quenched lime in the last phase of the autoclave. This is referred to as neutralization 501 in Figure 1. If this neutralization is not carried out, the resulting pulp will have a low pH, resulting in a significant amount of Fe in the solution as well as almost all copper.

An important feature of the process is that the pulp produced contains a minimum amount of iron in the solution (less than 100 ppm) and about 1-5 g / l copper in the solution. Choosing the right amount of slaked lime, fed to neutralization 501, these goals can be achieved even with concentrates that have a low metal-to-sulfur ratio and exhibit relatively high oxidation of sulfur, for example 15-30%. A typical example of such a concentrate and a petlandite / pyrite type mineral mixture of a mineral association.

However, for high metal concentrates: sulfur and / or low sulfur oxidation, the total amount of acid produced under oxidation under pressure 12 is less and no neutralization is required to produce a low pulp

content of iron and copper in the desired order. This embodiment of the method is referred to as Mode A and is described below with an explanation of Figure 4. A typical example of this type of concentrate is the petlandite / chalcopyrite type of mineral association.

In Mode A, the amount of acid consumed in the oxidation under pressure by other chemical reactions is more than sufficient to utilize / consume all the acid produced by the oxidation of sulfur.

Examples of the two modes A and C required for two different concentrates are shown in the table below:

type of chemical composition ratio% oxidation metal: sulfur sulfur process: concentrate

Si It is not Co S mode A 6.3 14 0.6 34 0.61 6 mode C 0.1 22 0.6 22 0.78 15

Thus, the first concentrate with Ni content of 14% exhibits only 6% sulfur oxidation during the pressure oxidation process and is therefore treated in mode A, whereas the second concentrate requires the use of mode C due to the higher oxidation (15%).

Mode C will be described with the explanations in FIG. 1.

Initially, the ore or concentrate is subjected to pressure oxidation 12 in an autoclave in the presence of a acid solution containing sulphate, chloride and copper ions. In the example shown, the amount of H ₂ SO ₄ introduced into the autoclave is about 40 g / l and the chloride concentration in the solution is about 10-12 g / l. Typically, the temperature is about 90 ° C to about 160 ° C at an oxygen partial pressure of 200-2,000 kPa. The retention time is about 0.5-5.0 hours and is inversely related to the temperature, usually the process is carried out continuously in an autoclave. However, if desired, the process can be run in series.

Neutralization 501 is effected by pumping lime (with a water content of 10-20% solid) in the last one or two autoclave chambers.

After pressure oxidation 12, the pulp obtained in the autoclave is drained through one or more expansion vessels 22 to reduce the pressure to atmospheric pressure and temperature to 90-100 ° C.

The slurry is then further cooled and filtered 24 to give a filtrate of oxidation under pressure 29 and a solid residue (filter cake of oxidation under pressure).

The neutralization step 501 is applied to obtain soluble copper precipitation on the filter cake from pressure oxidation, which would otherwise remain in the filter of the pressure oxidation product 29. Thus, neutralization 501 can be used to reduce to a minimum, the content of copper in the filtrate 29, which is usually less than 1 to 5 g / l of copper, facilitating the subsequent removal of copper from the solution. In addition, neutralization helps to reduce the iron content of the filtrate from oxidation under pressure. However, when slaked lime is added, it is preferable not to add too much, which would cause Ni / Co to precipitate. Typically, slaked lime is fed in such a way that the filtrate 29 of the oxidation under pressure is at a pH of between 3 and 4, which is considered appropriate for removing most of the copper with still minimal Ni / Co precipitation.

and the filtrate 29 from oxidation under pressure is usually subjected to recovery by dissolving 50 of copper, especially if there is a significant copper content in the starting concentrate so as to extract the copper contained and reduce as much as possible / Cu ²⁺ / in the refinery 63, typically below 100 ppm. In addition, the filter coke of oxidation under pressure is subjected to atmospheric leaching 14 to extract copper into the solution, which solution is subjected to extraction by dissolving copper 16. The leaching 14 is carried out with the purified product 120 from extraction of copper by dissolution 16, which product is a dilute acid of about 3 -20 g / l H ₂ SO ₄ . In addition, leaching 14 helps flush out the entrained solution containing a certain amount of Ni / Co outside the filter cake from the oxidation under pressure. These amounts, which will accumulate in stream 51, can be recovered at the main drainage (e.g. 1 to 10% of the stream depending on concentration) by settling at pH 7 to 8 with quenched lime in the form of Ni and hydroxides. Co, similar to the precipitate conditions 506, which is described below. Then the mixture of Ni / Co hydroxides can be filtered and recycled in the purification step 500, which is described below.

The slurry 31 obtained on leaching 14 is difficult to filter and the separation is solid (liquid is carried out by means of a series of thickeners according to flow decantation scheme 34 (PDD). The wash water is provided by a portion of the purified product from the extraction by dissolving 16, which is recovered at 36 and neutralized at 46 using a boil to remove the acid. The neutralization slurry 46 is filtered at 48 and gypsum waste is obtained and the liquid 51 is recycled as washing water.

with t

The charged wet extractant extractor 50 and 16 are separated 44 and then removed for electrolytic deposition of 20 on copper.

The extraction of copper by dissolving 50 and 16 is carried out with a common extractor. This is shown in Figure 2. Where the broken line means an organic extractant that circulates after separation 44. The separation 44 is carried out by consuming acid or electrolyte 55 from electrolyte deposition 20 to obtain a pure copper sulfate solution or rich / charged / electrolyte 57 which then undergoes an electrolytic deposition step 20. Any suitable copper extractant may be used that can selectively separate copper from the solution, which also contains Ni, Co.Fe and Zn. Hydroxyoxime such as LIX 84 ™ or UX 864 ™ reagents from Henkel Corporation is designated as the appropriate extractant.

If insignificant quantities of copper are present in the ore or concentrate, it is still advantageous to carry out the oxidation under pressure 12 in the presence of copper ions (eg 5 to 10 g / l Cu). The copper ions can be added in the form of copper salts such as CuSO ₄ or CuCI ₂ . Extraction is then carried out by dissolution and separation, but the electrolyte deposition 20 is omitted and the copper-rich liquid obtained from the dilution 44 of the organic extractants will be recycled into the pressure oxidation 12. Optionally, a copper concentrate may be added, in which case copper can be recycled after wet copper extraction and copper removal or electrolyte deposition removal for copper extraction. This will also be the case if lateritic ore is being processed.

Μ

The raffinate 63 is subjected to a purification operation

500 to obtain a Ni / Co solution free of elements such as Fe, Zn and Cu, which would cause difficulties in the subsequent steps of the wet extraction and electrolytic deposition of Ni and Co. Purification step 500 is a precipitation step in which residual Fe, Cu and Zn are precipitated by the addition of slaked lime and recycled MD / OH / ₂ . Typically, the solution fed to the purification step 500 will contain iron as well as some zinc and magnesium present in the concentrate. Sediment 500 is carried out at a pH of about 5 to 6, so ideally no more than 1 ppm Zn, 1 ppm Cu and 1 ppm Fe remain in solution. It is also important not to precipitate too much Ni / Co. This is achieved by strict pH control, ie. not allowing the pH to rise much. From this point of view, it is considered advantageous to recycle MD / OH / ₂ .

The precipitating product 500 is subjected to hard / liquid separation 502. Copper, iron and zinc, which are precipitated as hydroxides, can be returned to the process by a dilute acid solution, leaching solution 503, more specifically for the extraction of Ni / Co . The acid flushing product 503 is subjected to solid / liquid separation 505 to give mainly copper, iron and zinc hydroxides, which ensures the separation of zinc from the system. The liquid phase 504 of the solid / liquid separation 505 is recycled to the oxidation under pressure 12.

If the zinc content is high enough, Cu.Fe and Zn hydroxides can then be leached with a dilute acidic glass solution to separate zinc extraction. As a last resort, a wet zinc extraction step may be included if necessary.

The concentrations of Ni.Co and Mg in the solution after precipitation 500 will depend on the composition of the concentrate. Depending on the mineralogical composition, most of the magnesium in the concentrate may be leached during the oxidation under pressure.

12. For example, for Ni / Co concentrates containing, for example, 20% nickel and 5% magnesium, the solution after precipitation 500 will typically contain about 30 g / l nickel and about 6 g / l magnesium. When processing laterite ore, the magnesium content will be higher.

The solution obtained from the solid / liquid separation 502 is subjected to selective precipitation 506, whereby nickel and cobalt are precipitated as hydroxides or carbonates with suitable neutralizing agents such as slaked lime (Ca / OH / g), sodium carbonate, ammonia and sodium hydroxide. (Na (OH)). Neutralization and precipitation are carried out at a pH of 7 to 8, which results in minimal precipitation of Mg (OH) ₂ . The preferred neutralizing reagent is quenched lime as it is relatively the cheapest and the reaction does not introduce into the solution any other cations such as Na ⁺ n NH ₄ ⁺ .

Neutralization with slaked lime

NiSO _{4 (BQn)} + Ca (OH) ₂ -> Ni (OH) _{2 {TB)} + CaSO ₄ .2 H ₂ O _{TV) / 1 / gypsum

A similar reaction was carried out with CoSO ₄ to give Co / OH / ₂ and Mg / OH / _{2 respectively} .

Sodium base neutralization (NaOH)

NiSO _{4 (BOfl)} + NaOH -> Ni (0 ^ 2 (^ + _{NaSO 4} ( _BQn )! 2J

However, it is important to have a certain amount of Mg in the precipitated solid which facilitates the separation of Ni and Co as described below. It is appropriate to carry out two-stage countercurrent precipitation.

In some cases, precipitate with sodium hydroxide or ammonia is preferable, for example, to give a solid product (gypsum) so that the nickel precipitate is of higher quality and is calcium free.

The product of the sedimentation operation 506 is subjected to solid / liquid separation 508.

The solid / liquid separation fluid 508 is subjected to operation 510, preferably again with slaked lime, for the same reasons as above, to further precipitate Mg if necessary and thus to prevent the accumulation of Mg in system. The precipitation product 510 is subjected to solid / liquid separation 512. The solid phase of separation 512 is magnesium hydroxyl product 514. As mentioned above, part of the magnesium hydroxyl product 514 is recycled and used for precipitation 500. The liquid from separation 512 is is recycled in the pressure oxidation step 12 as indicated by the recycling stream 516.

A solid hydroxide cake from the separation step 508 containing a certain amount of Ni 0 Co is leached 518 with ammonia solution at a pH of about 6 to 8.

The ammonia solution may be ammonium sulphate or ammonium carbonate, but the former is considered to be better as it has a lower pH, which allows better separation in the Ni and Co solution. In addition, ammonium sulfate has a lower ammonium (gas) vapor pressure, as well as Ni and

Co is higher with ammonium sulfate. In this example, a 200g / l solution of ammonium sulfate is used.

During leaching to form nickel and cobalt soluble diaminesulfates, the following reactions occur:

(NH ₄ ) ₂ SO ₄ + Ni (OH) ₂ ^ Ni (NH ₃ ) ₂ SO ₄ +2 H ₂ O (3) (NH ₄ ) ₂ SO ₄ + Co (OH) ₂ -> Co (NH ₃ ) ₂ SO ₄ + 2 N ₂ O (4)

The available Mg in the solid phase is also dissolved as follows:

(Ni ₄ ) ₂ SO ₄ + Mg (OH) ₂ -> MgSO ₄ _ »+ 2 N ₂ O + 2NH ₃ (5)

During leaching, 518 does not attempt to leach 100% of the Ni / Co contained in the solid material, but only 90-99%. This allows leaching 518 to be carried out at low pH instead of at a high pH of about 9, which would otherwise be required. This higher pH requires the addition of ammonia to leach as a second reagent, together with ammonium sulfate.

A further problem that arises is that the extraction or commercially available Co extractant does not work effectively at this high pH value. The extractant is broken down and is not selective for Ni. As a result, it is first necessary to remove Ni instead of recovering Co, and then to reduce the pH by adding other reagents, such as acid, which means the preparation of ammonium sulfate by-product and the consumption of ammonia reagent. Another problem that arises is that in order to first wet the Ni extraction, it is necessary to oxidize the entire amount of Co to ⁺³ to avoid the extraction of Co with Ni. This oxidation is difficult to quantify. This further complicates the process. It is also necessary to reduce Co ³⁺ to Co ²⁺ with subsequent extraction of Ni and this is also difficult to perform.

In order to avoid the above difficulties, the process of the present invention provides for leaching 518 at a pH of from about 6 to about 8, and then the resulting solid fraction is further washed 520 with dilute ammonium sulfate solution as described below.

Another aspect of the method is that the concentration of Ni-ions in the solution during leaching 518 is controlled to remain at relatively low values, up to a maximum of about 10 g / l. As a result, a better recovery of Ni is obtained by leaching 518. With an established amount of Ni in the solid phase, the required amount of fluid can be calculated which would have to be supplied to obtain the desired Ni concentration. .

The leaching product 518 is subjected to liquid / solid separation 522.

Separation fluid 522 is co-extracted wet 534 to obtain Co-loaded extractant and raffinate, which is then subjected to wet-extracted Mg 536 to obtain Mg-loaded extractant and raffinate wet extraction of 538 Ni to obtain Ni-loaded extractant and raffinate. The raffinate obtained from the wet leaching 538 of Ni is recycled into leach 518.

The liquid / solid separation product 522 is recycled or washed 520 as indicated above, where the product is washed with ammonium sulphate solution. It is a weak ammonium sulphate solution with a concentration of about 10% and it is a leaching solution 518. It is obtained by washing the entrained solid ammonium sulphate solution 520.

Recycling product 520 is subjected to liquid / solid separation 524 and the solid is washed with water. The flushing water and the liquid / solid separation 524 are wetted to Co 526 extraction to obtain a Co-loaded extractant and raffinate to 527 Mg wet extraction to obtain the extractant. with Mg and raffinate, which is subjected to a wet 528 extraction of Ni to give an extractant loaded with Ni and a final raffinate, which is recycled at the recycling step 520.

In order to compensate for the water added to the water washing by filtration 524, a portion of the final purified solution is removed into concentrated ammonium sulfite raffinate coming from the wet extraction 538 of Ni. For this purpose, the concentrated ammonium sulfate cycle includes evaporation step 539 to compensate for the deviated raffinate from the dilute ammonium sulfate raffinate.

The wet extraction of cobalt 534,526, the wet extraction of Mg 536,527, and the wet extraction of Ni 538,528 were carried out with a common extractant, as in the wet extraction of Cu 50.16.

The extractant, which is considered suitable for both Co and Pb extraction, is an organic phosphoric acid extractant, in particular an organic phosphonium acid extractant such as Swapech 272 ™ w ^ Cyanamid Inc., which includes bis 2,4 , 4-trimethylpentyl phosphonic acid. Ni extraction is considered a suitable hydroxy-oxime based extractant, such as Henkel Corp.'s LIX 84 ™.

The charged Co, Ni, and Mg extractants are washed with suitable aqueous solutions to separate entrained ammonium sulfate solution and then dilute with acid to obtain pure Co and Ni-rich solutions and Mg-rich. liquid containing small amounts of Co and Ni Co and Ni solutions are removed for electrolytic deposition (EC) of Co and Ni in steps 530 and 532.

Prior to separation, the Co-loaded extractant is washed with a concentrated Co solution that deviates from the Co-rich solution taken to the electrolyte deposition / EC / of Co and / or a concentrated Mg solution which is separated from the Mg-rich liquid . This is done to facilitate the separation of Ni, which may be present in the extractant charged with Co. Similarly, the Mg-loaded extractant can be washed with a concentrated Mg solution, which is separated from a magnesium-rich liquid.

For the good separation of Co from Ni during wet Co extraction and wet Ni extraction, it has been found useful to have a certain amount of Mg in the solution fed to the Co wet extraction. Usually the solutions have the same Co: Ni ratio as the initial concentrate (usually 1:30). Thus for 10 g / l Ni corresponds to O.sub.3 / l Co.

The same extractant 534 and 536 is used for wet extraction of Co and Mg. The extractant is more selective for Co than Mg and more selective for Mg than Ni. In the wet 534 Co extraction, the amount of extractant used is limited to taking up all the available locations where there are more Co ions and Mg ions to a lesser extent, which reduces Ni extraction. In wet extraction &

of the Mg 536 accessible sites are filled mainly with Mg ions, to a lesser extent with Co ions and also a possible small amount of Ni ions. Ni and Co ions are recovered by recycling the Mg-rich fluid to precipitate Ni / Co 506 as indicated by arrow 543.

It has also been found advantageous to maintain the Mg concentration equal to the Co concentration, although it can vary over a very wide range from 1: 5 to 5: 1.

The benefit of having Mg is:

(i) minimize the amount of Ni recovered during wet Co extraction, while allowing for (ii) a high Co recovery rate, i.e. 90% and (iii) a high ratio of Co to Ni in the Co product, i. Co: Ni> 1000: 1.

Without the presence of Mg, there can be some compromise in the liquid extraction of Co such as:

(i) a certain amount of Ni is recovered together with Co or (ii) the recovery of Co is incomplete or (iii) the ratio of Co to Ni in Co product is too low.

When extracting Mg, a certain amount of Co / m.p. 510% / may remain un extracted during wet Co extraction and may instead be extracted during Mg extraction. Wet Mg Extraction Products are:

a / a rich liquid from the separation process and containing a certain amount of Mg, Ni and Co. It is recycled and not wasted; and b / Mg raffinate is very low in Co, ie.

at about 1 ppm, which allows the subsequent wet extraction of Ni to produce a very good ratio of Ni to Co in the Ni-enriched liquid entering for electrolytic deposition of Ni. This produces very pure cathode Ni and cathode Co.

The solid product from the solid / liquid partition 524 was washed (540) with dilute acid to recover entrained Ni / Co, and the liquid from the melt was recycled into the precipitate 500. The solid residue from the liquid / solid partition 542 was removed.

The suitable leaching temperature 518 of Ni / Co and the wet leaching of Ni / Co have been found to be in the range of about 30 ° C-60 ° C, preferably about 40 ° C-50 ° C.

Figures 3A and 3B describe the extraction of precious metals such as gold and silver. This method involves treating the final waste stream 35 of FIG. 1.

Precious metals are not recovered in the pressure oxidation process 12 but remain in the solid residue 35 which remains after the atmospheric leaching step 14.

In order to facilitate the recovery of the precious metals, drainage 22 from step 12 of the oxidation under pressure 12 is carried out in two steps. The first stage is at a temperature slightly above the solidification point of elemental sulfur, i. 120 ° C to 130 ° C, respectively, at a vapor pressure of about 50-150 kPa. Preferably, the operation is performed in a continuous mode, with the time for the first stage of drainage from the gateway being about 10 to 30 minutes.

The second stage of draining is at atmospheric pressure and the temperature is about 90-100 ° C for about 10 minutes. This allows elemental sulfur, which is still & molten in the first stage of draining, to quickly go into a solid phase, such as stable orthorhombic crystalline phase. This mode of action facilitates the production of pure crystals of elemental sulfur, which is important for the extraction of precious metals from the residual leaching.

The residual leaching 35 obtained during the atmospheric leaching step 14 contains, in addition to the precious metals, also hematite, crystalline elemental sulfur, unreacted sulfides (pyrite) and some other by-products which may be obtained from the concentrate used, for example gypsum and ferrous hydroxides.

In the remainder 35, it is believed that gold remains intact in the process and thus remains almost in its natural state. However, in the process of pressure oxidation 12, silver is oxidized and is likely to be in the form of silver salts such as silver chloride or silver sulfate.

It has been found that conventional cyanide cannot extract gold from residue 35. This is thought to be due to the encapsulation of gold in mineral particles such as pyrite. However, by pressure oxidation, gold can be released from these minerals and this is referred to as "complete oxidation leaching". In order to perform such leaching without oxidizing elemental sulfur, which is also contained in the residue 35, the process involves the step of separating as much elemental sulfur as possible.

First, the two stages of draining give good quality of sulfur crystals. Secondly, the leach residue 35 is subjected to foam flotation 402 to obtain a flotation concentrate 404 rich in sulfur and waste 406 with reduced sulfur content. The waste 406 is subjected to a solid / liquid separation 408 to obtain a liquid XO which is returned to the flotation process 402 via the contact vessel 410 and the solid product 412 is sent for complete oxidation leaching 414.

Flotation concentrate 404 was filtered (416) and dried to low humidity in an oven 418. The product was then fed to step 420 for sulfur leaching with sulfur extractants. Any suitable sulfur extractant such as perchlorethylene (PCE) or kerosene may be used. In this example, hot PCE is used. The leaching slurry 420 is filtered 422 and the resulting liquid is cooled 424 to obtain crystalline sulfur S0 and then filtered (425). The cooled sulfur can be optionally purified (not shown) to remove impurities such as selenium and tellurium. The solid sulfur is dried in an oven 426 to give sulfur product 428. The liquid phase from filtration 435 is recycled to leach 420 with hot PCE.

The solid filtration residue 422 is dried in an oven 430. The resulting product, which is a low sulfur content residue 432, is removed to a total oxidation leach 414.

Through the condenser 434, the PCEs from cooling 424 and drying 426 and 430 are recycled to leach 420 with hot PCE.

A test was performed in which 100 g of atmospheric leach residue 14 containing 25.1% elemental sulfur (S °) and 3% sulfide was treated by flotation 402 and leach 420. The thus obtained 73.8 g of desulphurized residue / NaOH full oxidation leaching material 414 / contains 1 _S 0% S ° and 4.1% sulfide, i. 6% total sulfur.

The desulfurized residue contained 5.9% elemental sulfur (S °) in the original leach residue, i. 94.1% was recovered as pure elemental sulfur.

The complete oxidation leaching 414 is carried out at about 200-220 ° and about 200-2000 kPa oxygen partial pressure, sufficient to fully oxidize all sulfur of the metal compounds to the highest valence. Thus all sulfur and pyrite are oxidized to sulphate. The oxidation is carried out in an acidic medium obtained from the acid obtained in the process. If enough pyrite is present, the reaction is exothermic and the process temperature can usually be reached. Typically, about 10% of the total oxidizable sulfur will be sufficient with a normal percentage of solid in the feed pulp.

After complete oxidation extraction 414, the slurry is neutralized 437 at pH 2-3 with limestone and then subjected to liquid / solid separation 438 by countercurrent decantation (PDD) to obtain a solid product containing noble metals and liquid 13 , which may contain a base precious metal such as copper. The liquid 13 may be combined with the fluid (stream 33) going to the liquid extraction 16 of copper as indicated in FIG.

A portion of the neutralized stream 51 (Figure 1) of the refinery from the liquid copper extraction 16 is separated into 49 and the resulting stream 53 is partially used (80%) as washing water in liquid / solid separation 438 and partially recycled (20% ) in complete oxidation leach 414 as indicated in FIG. The extraction cycle of the precious metals of FIGS. 3A and B is indicated as block / 55 in FIG. 1.

Prior to cyanidation 444, the solid product of partition 438 may optionally be subjected to lime treatment 443 to facilitate the extraction of silver by cyanidation 444 by decomposition of the silver jarosite compounds formed during complete oxidative leaching 414.

After partitioning, 438 precious metals remain in the solid residue. Now that pyrite and other encapsulating minerals of the starting concentrate are decomposed, the precious metals are suitable for cyanide 444.

In the cyanidation step 444, the solids were leached with NaCN in an alkaline medium. To achieve this, the solids are brought to a pulp state using a cyanide solution to form a pulp with a solids content of 30-40%. Additionally, NaCN and slaked lime are fed to maintain the required NaCN concentration and about 0.2 to about 0.5% g8l NaCN at a pH of about 10. The process is carried out at ambient temperature for about 4-8 hours in continuous operation mode.

Gold and silver are presented in large quantities in the cyanide solution and are usually recovered by established methods in the pulp carbon cycle, where activated carbon is added to the cyanide pulp to absorb the precious metals without the need for filtration. Charged carbon, already rich in precious metals, is removed by sieving (415) and the combined pulp is separated from the waste.

Charged charcoal is treated according to established methods for the extraction of precious metals. These are leaching, / electrolytic deposition / melting operations (447). The product is usually Dor metal, containing both gold and silver, which is sent for refining 449 of gold for the final separation of gold from silver.

2?

Purified barren carbon from coal regeneration 451 after the recovery of the precious metals is recycled into cycle 444 of the coal in the pulp.

The total extraction of precious metals in the whole process is usually well above 90% and under optimum conditions approaches 99%.

A test was performed in which the desulphurized waste was subjected to a complete oxidation leaching treatment at 220 ° C for 2 hours under pressure oxidation, after which pressure was removed and cooled to room temperature. The resulting pulp was neutralized to pH 3 with limestone and then filtered. The filter cake is then leached with cyanide solution under standard gold and silver extraction conditions.

After complete oxidation leaching 414 and cyanidation 444, the gold recovery is 97% at a NaCN consumption of only 1 kg / tonne. In comparison, the recovery of gold from the residue that was not oxidized by complete oxidative leaching 414 is only 34% and the cyanide consumption is extremely high - 19 kg / tonne NaCN.

Figure 4 is a flow chart of Mode A. The steps corresponding to those of the embodiment of Figure 1 are denoted by the same numbers.

The method involves a pressure oxidation step 12 in which the sulfide minerals of the concentrate or ore are oxidized at high oxygen pressure. Liquid / solid separation is then carried out. filtration / 24 to give a solid / filter cake from oxidation under pressure / 25 and filtrate 29 from oxidation under pressure.

The solid product 25 containing all or substantially all of the copper present in the feed concentrate,

2β is processed to recover copper 14 by acid leaching, wet extraction and electrolytic deposition as in FIG. 1. Thus, high quality cathode copper and a residue 35 containing noble metals are obtained. The residue 35 may be processed to recover the precious metals as described in the designations in Figs. 3A and B below. In Figure 4 this is denoted as block / 55.

The filtrate 29 was purified in 500 to remove harmful elements such as Cu, Fe and Zn by neutralization with slaked lime to pH 6, as noted in Fig. 1. Filtration yields a purified solution 36 containing Ni, Co and some other elements such as Mg that may be present in the feed solution.

The solution 36 is processed to extract Ni / Co as described by the indications in Figure 1. This is referred to as block 38 in Figure 4. The solid product 39 obtained from 38 is returned for pressure oxidation 12 to close the cycle as before (stream 516 in FIG. 1).

Although only preferred embodiments of the invention have been described in detail herein, the invention is not limited thereto and modifications may be made within the scope of the claims.

Claims (75)

1. A method of recovering Ni / Co from ore or concentrate, characterized in that the ore or concentrate is subjected to oxidation under pressure in the presence of oxygen and an acid solution containing halogen, copper and sulfate ions to form a liquid, containing Ni / Co from a pressure oxidation slurry, 'the liquid is selectively precipitated to give a solid product containing Ni / Co hydroxide; and the solid is subjected to Ni / Co leaching with ammonium solution to obtain a leaching solution containing Ni / Co and a residue. A method according to claim 1, characterized in that the residue is subjected to an acid wash to obtain a wash solution containing Ni / Co and a removable residue; and the wash solution is recycled to selective precipitation. A method according to claim 1, characterized in that the residue is subjected to an acid wash to obtain a wash solution containing Ni / Co and a residual residue; and the washing solution was treated to extract Ni / Co. Method according to any one of the preceding claims, characterized in that the selective precipitation consists of: the liquid is precipitated at a pH of about 5 to 6 for the precipitation of the iron and copper and zinc present in the liquid; and the resulting solution was precipitated at pH 7 to 8 to give a solid product containing Ni / Co hydroxide. 5. A method according to any one of the preceding claims, characterized in that the liquid containing Ni / Co is obtained by neutralizing the pulp of oxidation under reduced pH, wherein the pulp of copper, iron or zinc is present in the slurry. solid state, and Ni / Co are in solution. Method according to any one of Claims 1-4, characterized in that the liquid containing Ni / Co is obtained by pulping the oxidation under pressure to neutralize at a certain pH, wherein Ni / Co is in a solid state and subjected to acid leaching to give a Ni / Co solution. The method according to claim 2 or 3, characterized in that before the acid washing, the residue is flushed to obtain a second washing solution containing Ni / Co and an acid washed residue. The method according to claim 7, characterized in that one or both leaching solutions and the subsequent washing solution containing Ni / Co are subjected to wet Ni / Co leaching. A method according to claim 8, characterized in that wet extraction with nickel extractant is carried out to obtain a solution containing nickel. A method according to claim 8, characterized in that the wet extraction is carried out with a cobalt extractant to obtain a solution containing cobalt. A method according to claim 8, characterized in that the wet extraction consists of: ι wet extraction of cobalt in the presence of magnesium ions, using a cobalt extractant to obtain an extractant charged with cobalt ions and the first raffinate contains nickel and magnesium ions in solution; wet extraction of magnesium from the first raffinate using a magnesium extractant to obtain a magnesium extractant charged with magnesium and cobalt ions and a second raffinate; and wet extraction of nickel from the second raffinate is carried out using a nickel extractant to obtain an extractant charged with nickel and a third raffinate. Method according to claim 11, characterized in that each leach solution and subsequent leaching solution are wet-extracted, the third raffinate of | the wet leaching of the leachate solution is recycled to Ni / Co leaching and the third wet raffinate extraction of the next wash solution is recycled to the wash before the acid wash step. A method according to claim 12, characterized in that the wet extraction of Co, Mg and Ni is carried out with common extractants for Co, Mg and Ni. A method according to claim 11, characterized in that it comprises a step of separating cobalt and nickel from the dredged extractants to obtain cobalt and nickel solutions, respectively. 15. The method of claim 14, wherein, prior to the separation step, the cobalt-loaded extractant is washed with a concentrated cobalt solution, which is recycled from the dilution step. A method according to claim 11, characterized in that the leaching step of Ni / Co and the wet extracted M / Co are carried out at a temperature of about 30 ° C to about 60 ° C. A method according to claim 16, characterized in that the leaching step of Ni / Co and the wet extraction of Ni / Co are carried out at a temperature of about 45 ° C to about 50 ° C. A method according to claim 4, characterized in that it comprises the step of subjecting the nickel and cobalt electrolyte deposition solutions to the extraction of cobalt and nickel. 19. The method of claim 11, further comprising: dilution of the magnesium extractant to obtain a rich solution containing magnesium and cobalt ions; and the enriched solution is recycled at the stage of selective precipitation. 20. The method of claim 1, wherein, prior to dilution of the magnesium extractant, it is washed with a magnesium concentrated solution, which is a rich solution that is recycled from the dilution step of the magnesium extractant. I 21. The method of claim 11, wherein the cobalt extractant is similar to magnesium I extractant, the extractant being more selective for cobalt ί than magnesium. ι A method according to claim 8, characterized in that a cobalt solution and a first raffinate are obtained; and wet extraction of nickel from the first raffinate at a pH substantially the same as for wet extraction of cobalt to form a nickel solution and a second raffinate. A method according to any one of the preceding claims, further comprising the step of controlling the concentration of nickel in the leaching solution to a maximum value of 3 to 25 g / l. A method according to claim 23, wherein the maximum value is from about 8 to 15 g / l. The method of claim 24, wherein the maximum value is about 10 g / l. A method according to any one of the preceding claims, characterized in that selective precipitation is carried out in the presence of magnesium ions to produce a solid product containing magnesium together with Ni / Co hydroxide and a solution containing magnesium. A method according to claim 26, characterized in that it involves precipitation of the resulting solution to obtain a precipitate of magnesium hydroxide and a magnesium depleted raffinate. 28. The method of claim 27, comprising the step of recycling the depleted magnesium solution in the oxidation under pressure. 29. The method of claim 27 or 28, wherein at least a portion of the magnesium precipitate is recycled into the selective precipitation. A method according to any one of the preceding claims, characterized in that the leaching of Ni / Co is carried out with a solution of ammonium sulfate. 31. The method of claim 30, wherein the amine sulfate solution is at a concentration of about 150 to 250 g / l. The process of claim 31, wherein the ammonium sulfate solution is at a concentration of about 200 g / l. 33. A method according to any one of claims 1- 29, characterized in that the leaching of Ni / Co is carried out with a solution of ammonium carbonate. A method according to any one of claims 1-29, characterized in that the leaching of Ni / Co is carried out with a solution of ammonium sulfate and ammonium carbonate. A method according to any one of the preceding claims, characterized in that a copper salt or concentrate is added to the acid solution during the oxidation under pressure to provide copper in the acid solution. 36. The method of claim 35, wherein the copper is recycled to oxidation under pressure. A method according to any one of the preceding claims, characterized in that it comprises a step in which the liquid from the oxidation under pressure is subjected to wet copper extraction prior to selective precipitation. The method of claim 1, wherein the halide is chloride. A method according to any one of claims 1 to 34, characterized in that the ore or concentrate is a sulfide concentrate, which also contains copper and includes leaching steps of the pulp or solid product obtained by oxidation with acidic sulfate a leaching fluid comprising a solution of copper sulphate and a leaching residue; the leaching fluid is subjected to wet copper extraction to obtain a concentrated copper solution and depleted in copper raffinate; and at least part of the depletion of copper raffinate is recycled during the acid-sulphate leaching step. A method according to claim 39, characterized in that the liquid is subjected to wet copper extraction prior to selective precipitation to obtain a concentrated copper solution. A method according to claim 40, characterized in that it comprises a step in which the concentrated copper solution from wet copper extraction is subjected to electrolytic deposition to form metallic copper. A method according to any one of the preceding claims, characterized in that the ore or concentrate is a sulfide concentrate which also contains noble metals and includes the following steps: the elemental sulfur is removed from the pulp or solid product obtained by the oxidation under pressure to give a low sulfur residue; and the low sulfur residue is oxidized at high temperature and pressure to oxidize the sulfur, and from the precious metal compounds in the low sulfur residue a residual metal is recovered. 43. The method of claim 42, wherein the sulfur removal comprises: the pulp or solid product obtained from the oxidation under pressure is subjected to foam flotation to obtain a rich sulfide flotation concentrate and a sulfur-poor flotation waste; and the solid residue is subjected to sulfur recovery with a suitable extractant to obtain a low sulfur residue. 44. The method of claim 43, wherein the sulfur-poor flotation waste is subjected to a solid / liquid separation to produce a liquid that is recycled in foam flotation and a solid product that is subjected to oxidative leaching. 45. The method of claim 43 or claim 44, wherein the sulfur recovery is carried out at a temperature of 90-150 ° C. A method according to claim 45, wherein the sulfur extractant is selected from the group consisting of kerosene and perchlorethylene. A method according to claim 46, characterized in that the environmental leaching is carried out in an acidic medium at a temperature of about 200-220 ° C and an oxygen partial pressure of 500-1,200kPa. 48. A method of recovering Ni / Co from concentrates containing Ni / Co hydroxides, characterized in that: The Ni / Co concentrate was leached with ammonium solution to obtain a leaching solution containing Ni / Co and a residue; control the concentration of nickel in the leaching solution to a maximum of 3 to 25 g / l. The method of claim 48, wherein the maximum value is 8 to 15 g / l. 50. The method of claim 49, wherein the maximum value is about 10 g / l. A method according to any one of claims 48 to 50, characterized in that the leaching of Ni / Co is carried out with a solution of ammonium sulfate. 52. The method of claim 51, wherein the ammonium sulfate solution is at a concentration of about 150 to 250 g / l. A method according to claim 52, wherein the ammonium sulfate solution is at a concentration of about 200 g / l. A method according to any one of claims 48 to 50, characterized in that the Ni / Co leaching is carried out with ammonium carbonate solution. A method according to any one of claims 48 to 50, wherein the Ni / Co leaching is carried out with a solution of ammonium sulfate and ammonium carbonate. A method according to any one of claims 48 to 55, characterized in that: the waste is acid-washed to give a washing solution containing Ni / Co and a removal residue; the washing solution was selectively precipitated to give a solid product containing Ni / Co hydroxide; and the solid product is recycled to Ni / Co leachate. A method according to any one of claims 48 to 55, characterized in that: the residue is acid washed to give a wash solution containing Ni / Co and a removable residue; and the washing solution was treated to extract Ni / Co. 58. The method according to claim 56 or 57, characterized in that before washing the acid, a flushing of the residue is included to obtain a second flushing solution containing Ni / Co and an acid washable residue. 59. The method of claim 58, further comprising subjecting one or both of the leaching solutions and a subsequent washing solution containing Ni / Co to wet leaching to extract Ni / Co. A method according to claim 59, characterized in that the wet extraction is carried out with a nickel extractant to form a solution containing nickel. A method according to claim 59, characterized in that the wet extraction is carried out with a cobalt extractant to obtain a solution containing cobalt. 62. The method of claim 58, wherein: in wet cobalt extraction in the presence of magnesium ions, using a cobalt extractant to obtain a cobalt extractant loaded with cobalt ions and a first raffinate containing dissolved nickel and magnesium ions; wet extraction of nickel from the second raffinate is performed using a nickel extractant to obtain a nickel-charged extractant and a third raffinate. 63. The method of claim 62, comprising diluting cobalt and nickel charged extractants to obtain cobalt and nickel solutions, respectively. 64. The method of claim 63, wherein the solutions of nickel and cobalt are subjected to electrolytic deposition to extract nickel and cobalt. 65. The method of claim 63, wherein the magnesium extractant is diluted to obtain a rich solution containing magnesium and cobalt ions; and the rich solution is recycled to selective precipitation. 66. The method of claim 62, wherein the cobalt extractant is the same as the magnesium extractant, the extractant being more selective for cobalt than magnesium. 67. The method of claim 59, wherein: wet cobalt extraction is carried out at a pH of about 6 to 8 to give a cobalt solution and a first raffinate; and wet extraction of nickel from the first raffinate is carried out at substantially the same pH as cobalt wet extraction to obtain a nickel solution and a second raffinate. 68. Method for extracting Ni / Co from a solution characterized in that wet cobalt extraction is carried out in the presence of magnesium ions using a cobalt extractant to obtain a cobalt extractant loaded with cobalt ions and first raffinate containing dissolved nickel and magnesium ions; wet extraction of magnesium from the first raffinate is carried out using a magnesium extractant to obtain a magnesium extractant charged with magnesium and cobalt ions and a second raffinate; and wet extraction of nickel from the second raffinate is carried out using a nickel extractant to obtain an extractant charged with nickel and a third raffinate. 69. A method according to claim 68, characterized in that it comprises diluting the nickel and cobalt-loaded extractants to obtain the nickel and cobalt solutions, respectively. 70. The method of claim 69, wherein before the dilution, the cobalt-loaded extractant is washed with a concentrated cobalt solution that is recycled from the dilution. 71. The method of claim 70, wherein the solutions of nickel and cobalt are subjected to electrolytic deposition to extract nickel and cobalt. 72. The method of claim 71, comprising diluting the magnesium extractant to obtain a rich solution containing magnesium and cobalt salts. 73. The method of claim 72, wherein, prior to diluting the magnesium extractant, it is washed with a concentrated magnesium solution, which is a rich solution that is recycled from the dilution of the magnesium extractant. 74. The method of claim 68, wherein the cobalt extractant is the same as the magnesium extractant, the extractant being more selective for cobalt than magnesium. 75. The method of claim 68, wherein the wet extraction comprises: wet cobalt extraction at pH about 6 to 8 to give cobalt solution and first raffinate; and wet extraction of nickel from the first raffinate at virtually the same pH as for wet extraction of cobalt, to give a nickel solution and a second raffinate.